This invention relates to processes for recovering copper from copper containing feeds.
A number of alternative process routes exist for the treatment of copper ores, particularly ores in which the copper is present as copper sulphides such as chalcocite Cu2S, covellite CuS, bornite Cu5FeS4 and enargite Cu3AsS4. These routes include in-situ leaching, heap leaching and agitated tank leaching. Each route can incorporate different methods of regenerating ferric ion, the principal leach agentxe2x80x94bacterial oxidation, gaseous oxygen oxidation or chemical oxidants. The choice of route is influenced by factors such as resource tonnage and grade, mineralogy of both copper minerals and gangue, climate and environmental considerations.
These alternative methods are discussed in more detail below:
In-Situ Leaching
This technique has major environmental draw backs because of the difficulty in containing the leach solutions within the ore zone. Other difficulties arise from variable permeability of the ore, inability to control the leach reaction and the high likelihood of quite low overall recoveries. In-situ leaching is generally only considered for high permeability, low grade material which cannot be processed by other means and for resources where the leaching agent is quickly neutralised by waste rock surrounding the ore zone.
Heap Leaching
Heap leaching is commonly used to treat chalcocite ores in Australia e.g. Mt Gordon, Girilambone, and in other properties worldwide. However the effectiveness of this technique is highly dependent on the nature of the ore being treated. For some ores using medium height heaps, long leach times would be required for moderate recoveries. The rate of leaching is less dependent on the actual mineral leach rate than on those factors which will promote or inhibit leachingxe2x80x94oxygen supply, heap permeability, liquor percolation and ore grade variability. Many heap leach mine sites, experience continuing difficulty in attaining satisfactory stability in these factors with the result that heap leach performance often falls well below expectations which seemed theoretically reasonable at the time of the performance of trial leach tests.
Furthermore, where the ore contains high levels of pyrite there is a likelihood of a reaction of pyrite with ferric ion leading to increased acid production and conditions unsuitable for bacterial leaching.
Agitated Tank Leaching
Agitated tank leaching has the potential to maximise recoveries under controlled leach conditions. Leach times for either ground ore or a concentrate can be reduced to several hours. Pyrite reactions can be monitored and parameters such as oxygen supply and temperature adjusted to ensure the reaction is controlled. The process involves leaching ground ore in the presence of an acidic leachant containing ferric ion. The ferric ion oxidises the copper to form soluble copper ions and is itself reduced to ferrous ions. The ferrous ion is then converted back to ferric ion for further copper mineral attack. The ferrous ion oxidation can occur by a number of methods. The leachant can be separated from the solids and ferric regenerated bacterially or by reaction with acid and finely dispersed oxygen (or air). Alternatively ferric can be regenerated directly in the slurry by injection of finely dispersed oxygen or air. This can occur in the leach tank or by removing a small stream and passing it through a side-stream air/oxygen injection device. The process can be closely controlled so that the scale up to a full size plant is well understood and accurate estimates of final recoveries and operating and capital costs can be made. However there is significant effort and expense involved in regenerating the leachant. Furthermore the rate of leaching can decrease as ferric ion is used up in the leaching step. With some minerals e.g. covellite and enargite the leach rate is more dependent on the Eh which is strongly influenced by the ferric/ferrous ratio. Highest leach recoveries are often obtained by having an elevated ferric/ferrous rates at the end of the leach. This is difficult to attain in an atmospheric leach as the rate of oxygen dispersion in the pulp is limited.
Thus there is a need for a process which is suitable for treatment of a range of copper ores which deals with the disadvantages of the conventional processes described above.
A method for extracting copper from a mineral feed containing copper sulphide mineral including the steps of:
(a) leaching the feed with an acidic solution containing ferrous and/or ferric ions at a temperature above ambient in the presence of oxygen at superatmospheric oxygen pressure to produce a leachant solution containing copper ions;
(b) selectively extracting copper ions from the leachant solution by solvent extraction to form an extract solution containing copper ions and an acidic raffinate containing ferrous, ferric and low levels of copper ions;
(c) recycling some of the raffinate to be included in the acidic solution of step (a); and
(d) recovering copper from the extract solution.
Suitably the feed does not include significant quantities of soluble or leachable chloride. It should include less than 5% by weight of chloride more preferably less than 1% by weight. The presence of significant quantities of chloride ions makes for a very aggressive environment and thus requires the use of expensive equipment. This in turn leads to increased costs. Hence the preference for feeds low in chloride. Similarly the acidic solution should be kept substantially free of chlorides.
The copper sulphide mineral may include chalcocite Cu2S, covellite CuS, bornite Cu5FeS4 or enargite Cu3AsS4 and/or mixtures of two or more of these. The copper sulphide mineral may include iron mineralization. The iron mineralization may be pyrite. It may include 15% to 95% pyrite. The copper sulphide mineral may be from the Esperanza deposit in Australia. Esperanza ore typically contains about 70% pyrite in association with chalcocite. After mining dilution the ore will average 60% pyrite i.e. the feed to the autoclave will be 30% to 70% pyrite averaging 60%.
By comparison an ore concentrate from the Mammoth deposit in Australia, after flotation may typically contain about 30% by weight of pyrite. This concentrate is also suitable for treatment to recover copper by the process of this invention.
To facilitate more rapid leaching the feed may be ground. After grinding the majority of particles in the feed have a size generally less than 150 microns more preferably less than 100 microns. Preferably after grinding the feed will be 80% by weight passing 150 microns and more preferably 80% by weight passing 75 to 106 microns.
In step (a) the acidic solution will contain sulphuric acid. The sulphuric acid is suitably generated in-situ in step (a) by the oxidation of sulphides contained in the feed and by transfer via solvent extraction from electrowinning. Suitably the concentration of the acidic solution falls within the range 10 to 60 g/l H2SO4.
Ferric ions in step (a) may be generated in-situ by the leaching of iron in the feed to produce ferrous ions and by the oxidation of ferrous ions in the feed and recycled raffinate to ferric ions by oxygen. The ferric ions promote the dissolution of the copper minerals to produce copper ions (mostly divalent cupric ions Cu2+) and in the process are themselves reduced to the ferrous Fe2+state. Thus the acidic solution will contain both ferric and ferrous ions. It is preferred that the ratio of ferric to ferrous is at least 1.0 and is more preferably in excess of 2.0 at the end of the autoclave stage. Suitably the concentration of iron in solution is maintained in the range 10 to 40 g/l during leaching.
Suitably the leaching step (a) is carried out at a temperature in the range 50xc2x0 C. to 105xc2x0 C., more preferably 65xc2x0 C. to 95xc2x0 C. The oxygen partial pressure used for carrying out step (a) is generally maintained in the range 1 to 10 Bar more suitably 2 to 7 Bar.
There may be more than one leaching step. Subsequent leaching steps may be carried out under similar conditions to step (a). Alternatively, one or more subsequent leaching steps may be carried out under normal atmospheric conditions. Where prior or subsequent steps are carried out under normal atmospheric conditions, it is important to ensure that there is sufficient excess ferric ion in solution from step (a) to continue substantial leaching. Under normal atmospheric conditions with air/oxygen sparging there is greatly reduced regeneration of ferric ion from ferrous ion. Generally speaking the leaching steps should result in at least 80% and more preferably 90% of all copper in the feed being extracted into the acid solution for leaching. The leaching in subsequent tanks is enhanced if waste oxygen gas from the autoclave step is sparged into these tanks.
The solvent extraction may be carried out in one or more steps. It may be carried out using an organic phase containing a specific copper ion extractant. The organic phase may include a diluent, usually a kerosene with 10-30% aromatic content and an elevated flash point. The diluent reduces the viscosity of the organic phase and enhances organic/aqueous phase separations. Extractants may include modified aldoximes or modified or unmodified aldoxime/ketoxime mixtures. The following summarises a range of contents of a number of possible reagents for solvent extraction.
Reagent 1
aldoxime: 5-nonylsalicylaldoxime
Modifier: 2,2,4-trimethyl-1,3-pentanediol diisobutyrate and high flash point diluent
Reagent 2
ketoxime: 2-hydroxy-5nonylacetophenone oxime
aldoxime: 5-nonylsalicylaldoxime
Modifier: Tridecanol and high flash point diluent
Reagent 3
aldoxime: 5-nonylsalicylaldoxime
Modifier: Tridecanol and high flash diluent
Reagent 4
ketoxime: 2-hydroxy-5nonylacetophenone oxime
aldoxime: 5-nonylsalicylaldoxime and high flash point diluent
Reagent 5
aldoxime: 5-nonylsalicylaldoxime
Modifier: Tridecanol and high flash point diluent
The pregnant liquor solution from leaching may typically contain 15 to 30 or 40 grams/litre of copper. About 25 g/l of copper is particularly appropriate. After contact with the organic phase in solvent extraction the depleted copper solution or raffinate will contain about 2-4 g/l copper. There is no need to reduce the copper tenor of this stream to lower levels as the majority or all of the raffinate is recycled to the head of the leach. The presence of copper in the recycle stream can also catalyse the regeneration of ferrous to ferric ions in the leach.
The loaded organic resulting from solvent extraction may be stripped of copper by contacting it with high strength acid (e.g. 180 g/l sulphuric acid) to produce an electrolyte stream. The electrolyte stream may be fed to an electrowin circuit to recover copper.
Spent electrolyte may then be recycled to recover more copper.
In a typical process for carrying out the invention, after leaching in an autoclave and following further leaching and cooling, the leached slurry may be discharged to a hydroclassifier. Typically no flocculent is added to the hydroclassifier. Underflow solids from the hydroclassifier may reach a final settled density of 55-65% solids. The underflow solids are passed to a storage tank before flocculation and filtration on a horizontal belt filter. The filter cake may be washed in three stages counter-currently on the residue filter belt. This removes copper present in the residual moisture of the filter cake.
The overflow from the hydroclassifier contains fine solids that did not settle in the hydroclassifier. It is preferred that some solids, typically 5,000-100,000 ppm, be contained in the hydroclassifier overflow.
Removal of the fine solids from the feed to the leach residue horizontal belt filter greatly increases the filtration rate of residue solids and allows for improved wash efficiency. A removal of about 3% of the solids by weight in the hydroclassifier overflow allows a substantial improvement in the filtration rate on the horizontal belt filter.
The presence of fine solids in the hydroclassifier overflow also improves the coagulation/flocculation properties of this stream. This allows sufficiently large flocs to be formed which can then be satisfactorily removed in the clarifier.
The hydroclassifier can be fitted with recycle of hydroclassifier overflow back to the hydroclassifier and increase the proportion of fine solids reporting to the overflow.
Correct selection of coagulants/flocculants for treatment of the fine solids in the hydroclassifier overflow can ensure almost total removal of fine solids from the pregnant liquor stream. Any fine solids present in the pregnant liquor stream can cause disruption in the subsequent solvent extraction stage by the formation of stable emulsions known as crud.
The coagulated and flocculated fine solids can be removed in a clarifier known as a pinned bed clarifier. The coagulated and flocculated solids are introduced to the clarifier through a centre feed-well. Large flocs settle to the bottom of the clarifier and are discharged through a bottom outlet. Any fine remaining solids can be carried upwards through a bed of fine polystryene beads about 1 mm in diameter. These beads can be held in place or xe2x80x9cpinnedxe2x80x9d by retaining means such as a retaining wedge wire screen. The bed of beads may be about 600 mm deep. The fine flocs are trapped on the beads and form a flocculated bed which can capture further fines carried up to the beads. The clarified liquor passes through the beads and screen and may be discharged to sand filters for final polishing before solvent extraction. Periodically liquor may be removed from the clarifier via a valve. This may be located near the base of the clarifier. This creates a back-flow through the bed and removes solids build-up. These solids can then be discharged through the bottom outlet.
The flocculated fine solids discharging from the bottom of the pinned bed clarifier may be further treated in a multistage counter current decantation (CCD) circuit. Wash water can be added to this circuit to wash soluble copper away from the near barren fine solids. Liquor discharging from this circuit can be further treated in a two stage extraction xe2x80x9cbleedxe2x80x9d solvent extraction circuit to remove copper to low levels. The xe2x80x9cbleedxe2x80x9d solvent extraction circuit can use barren organic diverted from the main solvent extraction circuit. Raffinate from the xe2x80x9cbleedxe2x80x9d solvent extraction circuit is combined with the washed CCD underflow solids and filter cake from the residue filter belt for neutralisation prior to disposal in the tailings dam.
Alternatively the fine solids discharging from the pinned bed clarifier underflow can be added to the hydroclassifier underflow solids and the combined stream filtered on the residue belt filter. Whilst the removal of fines from the hydroclassifier underflow greatly increases the filtration capacity of this stream there is little or no adverse impact on the filtration rate if those fines are correctly coagulated and flocculated and remixed with the coarse fraction. This method does not require a CCD circuit for the washing of the fine solids. Any required bleed from this circuit can be taken from the solvent extraction raffinate before final copper scavenging, neutralisation and disposal.
The applicant""s process has a number of advantages over the prior art. For instance, by being carried out in reaction vessels such as an autoclave it does not have the environmental problems associated with in-situ leaching. The extractive processes can also be more carefully controlled in reaction vessels than in in-situ processes not employing reaction vessels. Furthermore it does have the significant advantage that it involves in-situ generation of ferric ion in the reaction vessel. This avoids the need for a separate ferric ion regeneration step and also makes it possible to generate sufficient ferric ion in the first leaching step as to promote downstream leaching with the excess ferric ion generated in the first step. This avoids the need for a separate and costly step to regenerate ferric ion downstream while at the same time making use of the excess heat generated in the first step to promote downstream leaching reactions.
A preferred embodiment of the present invention will now be described with reference to the accompanying drawing.